The present invention relates to a process by which, in connection with an electrolytic zinc process and particularly the leaching process for zinc calcine, the recovery of lead, silver and gold from the iron-bearing residue is also effected in addition to a high recovery of zinc, copper and cadmium, in an economical and simple manner.
The starting material of an electrolytic zinc process is a sulfidic zinc concentrate, from which an oxidic product, zinc calcine, is obtained by roasting. This calcine contains, in addition to the principal constituent, zinc oxide, practically all of the iron of the original concentrate, combined with zinc as zinc ferrite. The iron content in the concentrate usually varies between 5 and 15%, depending on the concentrate. An iron content of about 10% in the concentrate represents a typical value of currently used raw materials. This means that about 10% of the zinc of the concentrate is bound in zinc ferrite, ZnFe.sub.2 O.sub.4, the content of which in this typical case is 21.5% of the total amount of calcine.
In addition to zinc, the zinc concentrate also contains other valuable metals such as Cu, Cd, Pb, Ag and Au, and the recovery of these metals is of considerable significance for the total economy of the zinc process. However, in planning a zinc process or in modifying a process, it is necessary to take into account the behavior of several elements present in the concentrate in the process. Some of these elements (Zn, S, Cu, Cd, Pb, Ag, Au) are of primary importance for the economy of the zinc process, whereas others (Fe, Cu, Ni, Ge, Tl, In, Ca, Mg, Mn, Cl, F) have less or no economic importance but have to be taken into account precisely, with regard to the functioning of the process. In addition, there are elements which are significant in terms of environmental protection (S, Hg, Se), the quality of the byproducts (Hg, Se, As, Sb, Sn), or waste formation (Fe, Si, Al, Ca).
It is of primary importance for the economy of the process that the recovery of zinc is high. In a process alternative which can be considered good at present, the target set for the recovery of zinc must be at minimum 97-98%, and also there must be a maximally good recovery of the above-mentioned valuable elements in a saleable form.
The following approximate values can be taken as average valuable-metal contents in a typical zinc concentrate: Zn 53%, Cu 0.5%, Cd 0.2%, Pb 1%, Ag 60 g/t, Au 0.5 g/t. This means that, at the current prices of the products, the total value content of copper and cadmium jointly is 4-5%, that of lead, silver and gold 8-10%, and, furthermore, the value of sulfur calculated as sulfuric acid 5-6%; i.e. the value content of the byproducts is approximately 20% of the value of zinc, which is the principal product of the process. Thus it is evident that a maximal recovery of the said byproducts is also essential for a competitive process.
As regards the said harmful elements, especially iron, its recovery does not have special economic importance (the value of the iron as iron ore is about 0.2% of the value of the zinc); instead, the iron compounds formed during the process often cause a waste problem difficult to solve.
Prior to 1965, it was common in an electrolytic zinc process to recover the zinc present primarily as zinc oxide and zinc sulfate by means of a dilute acid leach, whereas the undissolved ferritic material constituted a leach residue, which in several cases was directed to waste disposal areas. In such cases, zinc, copper and cadmium bound in the ferrite, as well as lead, silver and gold which remained in the form of insoluble compounds under the leaching conditions were also lost in the waste disposal area along with the iron detrimental to the process. At that time, the degrees of recovery of the metals were typically 87-89% for zinc, approx. 50% for copper, 50-60% for cadmium, and 0% for lead, silver and gold. The amount of ferritic leach residue was on the average approximately one-third of the amount of calcine fed into the process. The said procedure was applied, since a suitable method was not known for the separation of the large iron amounts present in the calcine.
An essential improvement in this respect was provided by the patent applications filed in 1965 by Steintveit and by Haigh & Pickering (Norwegian Pat. No. 108047 and Australian Pat. No. 401 724). In the processes disclosed in these patent applications, the ferrites were leached and the iron was precipitated in the form of well-settling and filtrable jarosite compound. In the former process, the iron was precipitated under atmospheric conditions by using the zinc oxide of the zinc calcine for the neutralization of the sulfuric acid produced during the precipitation. In the latter process, the iron was precipitated in an autoclave, without neutralization. The jarosite process as a process in accordance with the former patent, supplemented with an acid wash (Norwegian Pat. No. 123 248), has found extensive use in the zinc industry. The process is described, for example, in G. Steintveit's article "Die Eisenfallung als Jarosit und ihre Anwendung in der Nassmetallurgie des Zinks", Erzmetall 23 (1970) 532-539.
In the jarosite process, the yield of zinc rises to 97-98%, the yield of cadmium to 90-95%, the yield of copper to 80-90%, and the yields of lead, silver and gold to 70-80%. A jarosite precipitate, the iron content of which is approximately 30% an amount somewhat less than 30% of the amount of calcine fed into the process, is removed from the process. The precipitate often--especially owing to its high annual output--constitutes a waste problem for the industrial establishment concerned. A leach residue which contains most of the lead, silver and gold of the concentrate is removed from the leaching stage of the process. The amount of the leach residue is usually approximately 5% of the amount of the calcine feed. The lead content in the residue is usually approximately 20%. The low head content of such a leach residue and its oxidic and sulfatic composition have lowered its commercial value, and therefore it is understandable why earlier, at a time of a relatively low price level of lead and noble metals, it did not offer an especially interesting material for processing and was in many cases directed to the waste disposal area together with the jarosite precipitate.
Soon after the emergence of the jarosite process, Societe de la Vieille Montagne developed the goethite process (Belgian Pat. No. 724 214). It differs from the jarosite process as regards the iron reduction stage (Fe.sup.3+ .fwdarw.Fe.sup.2+) and the iron precipitation stage. The iron is precipitated as goethite by using the zinc oxide of the zinc calcine for the neutralization of the sulfuric acid produced during the precipitation.
The metal yields of the goethite process are in the main the same as those of the jarosite process. The iron precipitate and the leach residue are removed from the process. The latter is similar to the leach residue of the jarosite process in both quality and quantity. The iron precipitate is in this case goethite-based, and its iron content is approximately 45-48%. Its amount is clearly less than that of the corresponding precipitate in the jarosite process, but even in this case it is nearly 20% of the amount of the zinc calcine feed. The goethite process has been described in the article by J. N. Andre and N. J. J. Masson "The Goethite Process in Retreating Zinc Leaching Residues", AIME Annual Meeting, Chicago, February 1973.
As is evident from the above brief descriptions of the processes, both the jarosite and the goethite process produce relatively large amounts of iron precipitate, which is not suitable for, for example, the production of crude iron without further treatment, and for which no other use has been found, but the precipitates have as a rule been directed to waste disposal areas.
The attempt to diminish the waste problem has lead to a search for process alternatives in which the iron can be separated in the form of sufficiently pure hematite with the purpose of channeling it to the iron industry as raw material. On this basis, there have been developed the hematite processes, in which the iron is precipitated as hematite out from the process solution during an autoclave stage. The first hematite process was developed by The Dowa Mining Company, and the process is in use at a zinc plant in Iijima, Japan. The process has been described in the article by S. Tsunoda, J. Maeshiro, E. Emi, K. Sekine "The Construction and Operation of the Iijima Electrolytic Zinc Plant", TMS Paper Selection AIME A-73-65 (1973).
Another hematite process was recently developed by Ruhr-Zink GmbH in the Federal Republic of Germany. The process has been described in DT-OS 26 24 657 and DT-OS 26 24 658 and in the article by A. von Ropenack "Die Bedeutung der Eisenfallung fur die hydrometallurgische Zinkgewinnung, Erzmetall Bd 32 (1979) 272-276.
Outokumpu Oy has developed a process based on the utilization of jarosite compounds, i.e. the conversion process, in which special attention has been paid to a high recovery of zinc, copper and cadmium and to the simplification of the process for leaching the zinc calcine. The process has been in use at the Kokkola zinc plant of Outokumpu Oy since 1973. At the time that the process was adopted, the raw material of the plant was so low in lead, silver and gold that the recovery of these elements did not seem economically advisable at the then prevailing relative prices. On the other hand, it was viewed as advisable to aim at a maximally high recovery of the zinc, copper and cadmium present in the concentrate and at simplicity of the apparatus and the processing method. It was proven that by giving up the individual separation of the leach residue which contained lead and noble metals, normally carried out in connection with the previously described jarosite process, it was possible to combine the stages normally included in the jarosite process--ferrite leach, (preneutralization), jarosite precipitation and acid wash of jarosite precipitate--to form one stage in which the ferrite dissolves (consuming acid) and the iron simultaneously precipitates as jarosite (producing acid) and thereby to simplify the process of leaching zinc calcine. In this case the reactions (1) and (2) representing the phenomena occurring in the process
(1) 3ZnFe.sub.2 O.sub.4(s) +12H.sub.2 SO.sub.4(aq) .revreaction.3ZnSO.sub.4(aq) +3Fe.sub.2 (SO.sub.4).sub.3(aq) +12H.sub.2 O.sub.(aq) PA1 (2) 3Fe.sub.2 (SO.sub.4).sub.3(aq) +Na.sub.2 SO.sub.4(aq) +12H.sub.2 O.sub.(aq) .revreaction.2NaFe.sub.3 (SO.sub.4).sub.2 (OH).sub.(s) +6H.sub.2 SO.sub.4(aq) PA1 (3) 3ZnFe.sub.2 O.sub.4(s) +6H.sub.2 SO.sub.4(aq) +Na.sub.2 SO.sub.4(aq) .revreaction.2NaFe.sub.3 (SO.sub.4).sub.2 (OH).sub.6(s) +37ZnSO.sub.4(aq) PA1 (4) PbSO.sub.4(s) +Na.sub.2 S.sub.(aq) .revreaction.PbS.sub.(s) +Na.sub.2 SO.sub.4(aq) PA1 (5) 2AgCl.sub.(s) +Na.sub.2 S.sub.(aq) .revreaction.Ag.sub.2 S.sub.(s) +2NaCl.sub.(aq).
are in mutual interaction and form a sum reaction (3), in which the zinc of the zinc ferrite passes into the solution and the iron is converted during the same stage via the solution to the jarosite phase. The leach yields and total yields of zinc are respectively 98-99% and 97.5-98.5%, and the total yields of copper and cadmium are 85-90%. The process is described in Finnish Patent Application 410/73 and in the articles by T-L Huggare, S. Fugleberg, J. Rastas "How Outokumpu Conversion process raises zinc recovery", World Min. (1974) 36-42 and by J. Rastas, S. Fugleberg, L-G Bjorkqvist, R-L Gisler "Kinetik der Ferritlangung und Jarositfallung" Erzmetall Bd. 32 (1979) 117-125.
On one hand, as the raw material range has come to contain more lead, silver and gold than previously, and on the other hand, as the changes in the relative prices of these metals--especially those of the noble metals--nowadays make it necessary to plan the leach process of the zinc calcine so that, in addition to a high recovery of zinc, copper and cadmium, a corresonding recovery is achieved also regarding lead, silver and gold.
In the jarosite and goethite processes summarized above, there is produced during a hot acid leaching stage a leach residue which no longer contains ferrites but contains all of the lead, silver and gold contained in the calcine fed to the neutral leaching stage. The lead content of this leach residue is in general relatively low, often about 20%. The low lead content of the residue and its oxidic and sulfatic composition decrease its commercial value. Therefore, it is understandable that processes by which the lead, silver and gold can be obtained in a more saleable form have been developed for the further treatment of this leach residue--originally intended for sale.
Asturiana De Zinc S. A. has, in its Finnish patent application No. 3435/70, disclosed a process in which the leach residue of the hot acid leach, produced in the manner described above, the lead being present in the residue as lead sulfate and silver as silver chloride and silver sulfide, is leached by means of a chloride-saturated and acidified solution in the presence of compounds which accelerate the oxidation of the metal sulfides present in the residues, such as copper chlorides, at a temperature which is between the ambient temperature and the boiling point of the solution, the leaching taking place in one or several stages. Thereby, both silver chloride and lead sulfate dissolve, forming silver and lead chloride complexes. The conversion of silver sulfide to silver chloride is promoted by additions of suitable reagents such as copper chloride. Both lead and silver can be separated from the solution as insoluble salts, such as sulfides, or by precipitating the metals successively out from the solution by using lead and zinc as cementing reagents.
Finnish patent application No. 761582 of Societe des Mines et Fonderies de Zinc de la Vieille Montagne relates to a process in which, on one hand, noble metals, especially silver, and on the other hand, lead are recovered from the leach residue of a hot acid leach, the residue no longer containing ferrite. The process is characterized in that the leach residue is slurried in water, the pH of the slurry is adjusted to between 1 and 5, a sulfide collector agent is added, and the slurry is froth-flotated. The products are, on one hand, a sulfide concentrate which contains silver, sulfides--above all, silver sulfide and zinc sulfide--and elemental sulfur, and on the other hand, a froth-flotation residue. The pH of the froth-flotation residue slurry is adjusted to between 1 and 4, an organic anionic collector agent is added, and the slurry is froth-flotated. The products obtained are a lead sulfate concentrate and a froth-flotation residue, which contains silica, iron oxides and calcium sulfate.
Finnish patent application No. 214/74 of Asturiana de Zinc S. A. also relates to a process in which lead and silver are recovered by froth-flotation from a leach residue which has been obtained from a hot acid leach of the neutral leach residue in a zinc process and no longer contains ferrites. The process is characterized in that, at first, most of the silver, sulfur and zinc are froth-flotated using suitable collector agents without sulfuring agents, the procedure is repeated on the froth-flotated product 1-3 times, whereby a concentrate concentrated with regard to silver, sulfur and zinc is obtained, whereas the residue is treated with an agent which activates the surface of lead sulfate, preferably sodium sulfide, whereby the surface of the lead sulfide present in the residue is activated, and when a suitable collector agent is added, the lead sulfate thus masked is froth-flotated. The procedure is repeated on the obtained product 1-3 times, whereafter the final lead sulfate concentrate is obtained.
A Japanese company, Mitsubishi Metal Corporation, disigned as early as 1961 a process for the recovery of silver by froth-flotation from the ferritic leach residue of the leaching of zinc calcine. The ferritic leach residue is produced by treating the neutral leach residue under mildly acidic conditions (pH=1.8), whereby the free zinc oxide dissolves, and the zinc ferrite remains undissolved. The process is described briefly in the article by A. Moriyma, Y. Yamamoto "Akita Electrolytic Zinc Plant and Residue Treatment of Mitsubishi Metal Mining Company, Ltd.", AIME World Symposium on Mining & Metallurgy of Lead and Zinc, Vol. II, 1970, 198-222. The process has been described later--in the form it was at that time and in greater detail than described previously--in the article by Y. Yamamoto "Silver Recovery from Zinc Residue", TMS Paper Selection AIME A 77--18 (1977). According to the method, the ferritic leach residue is slurried in water, a sulfide collector agent and a frothing agent are added to the slurry, and the slurry is froth-flotated. Primarily the sulfidic phases of the ferritic leach residue, sphalerite (ZnS) and argentite (Ag.sub.2 S), rise into the froth. The yields of silver and gold by the method are respectively 75-80% and 30-35%.
An examination of the methods for the recovery of lead an silver implemented or suggested in connection with processes for leaching zinc calcine shows that the process alternatives disclosed in Finnish patent applications No. 3435/70, 761582 and 214/74 relate to the further treatment of a leach residue--which is a leach residue from a hot acid leach of a neutral leach residue--and that the methods are thus connected with the use of multiple-stage jarosite and goethite processes. When a further treatment of the residue from a hot acid leach, or often also a further treatment of a material called the strong acid leach residue, by means of either chloride leaching or froth-flotation is added to these processes, an entity consisting of several partial processes is obtained, and the difficulties encountered in the technical control of the process are in proportion to the complexity of the total process.
The process of Mitsubishi Metal Corporation relates to a ferritic leach residue and, within it, specifically to the recovery of silver. This recovery is carried out by means of a direct froth-flotation of the leach residue. The lead present in the leach residue cannot be recovered by the process, the recovery of gold is low, 30-35%, and the recovery of silver also remains between 75 and 80%.
If we examine the process according to Finnish patent application no. 410/73 of Outokumpu Oy, we see that the technical implementation of the leach process is simple in terms of both the apparatus and the process control. It has a deficiency in that the lead, silver and gold present in the concentrate cannot be recovered by means of it, and these elements pass to the waste disposal area along with the jarosite precipitate. The present process is an improvement which eliminates this deficiency of Finnish patent application No. 410/73.
It has now surprisingly been observed that lead, silver and gold can be sulfidized selectively from a ferritic residue without simultaneous sulfidization of the zinc present in the solution and without substantial simultaneous dissolving of ferrite.